Hollow rock bolts were designed to make the grouting process more reliable (Figure 33.11). In hollow rock bolts, there is a hole in the center of the rock bolt. This hole is used for the grouting process and entangling tubes are eliminated. Hollow rock bolts may have the same cone and wedge anchor mechanism.
When the rock bolt is downward facing, the grout is inserted through the hole at the center and a short tube is inserted outside of the bolt to remove the trapped air.
When the rock bolt is upward facing, a short tube (inserted outside of the bolt) is used for the grouting process and the central hole is used as a breathing tube.
Rock bolts are a type of drilled soil nail or anchor used when the ground to be stabilized consists mostly of rock materials. These types of inclusions are typically grouted in place and posttensioned, similar to soil anchors (Figure 15.8). The bolts hold potentially unstable jointed or fractured rock masses together in compression with bearing faceplates providing passive resistance, forming a more stable structural entity. Rock bolts have been used for many years as temporary roof support in the mining industry and for tunneling, and are now used routinely for stabilizing roadcuts, rock cliffs, steep slopes, bridge abutments, and reinforced concrete dams. Two other versions of rock bolts sometimes used are those that have expansive shells (like a drywall anchor) rather than grout for resistance, and untensioned steel rods grouted into boreholes referred to as dowels (Hausmann, 1990).
Anchors and rock bolts (Fig. 4.35) involve the same mechanical principle, but they are employed in soft soil and rock, respectively. Anchors and rock bolts are active reinforcing elements designed to anchor and stabilize the rock mass during tunnel excavations. In case of rock mass movement, the bolts will develop a restraining force that is transferred back to the rock mass as in Fig. 4.36. The driving force is countered in this way, such that the total resistance mobilized within the rock mass is equal to the applied force.
220.127.116.11 Lattice Girders
Lattice girders (Fig. 4.37) are lightweight triangular steel frames. In crown heading excavation, for example, immediate support should be applied to the excavation with lattice girders. They are used in conjunction with shotcrete and act as armor to achieve a good level of support in tunneling. To provide additional forward support, strata bolts can be inserted through the lattice girders (Tunnel Ausbau Technik, n.d.).
I beam girders (Fig. 4.38) are relative heavy steel frames. They are widely used in China tunneling projects for weak rock formation. However, the use of I beam girders will have a bad influence on shotcrete quality due to their geometric shape.
18.104.22.168 Sprayed Concrete
Sprayed concrete is a fast hardening material used to stabilize and support tunnels. It comes in two forms: gunite (dry mix) and shotcrete (wet mix). Shotcrete is more efficient than gunite as shotcrete has a lower rebound rate.
Dry sprayed concrete is used when small amounts and high early strength is needed (Höfler, Schlumpf, & Jahn, 2011). The material is conveyed in a dry or semidry state to the nozzle, where water is added to the mix before being applied at high velocity onto the substrate.
Shotcrete is used more when a high concrete quality is required (Höfler et al., 2011). It requires a mechanized nozzle, as the wet mixture is too heavy to be held by a worker (Fig. 4.39). This wet sprayable concrete is workable, premixed, and consists of aggregate, cement, and water. For spraying, wet-sprayed concrete is mixed with air and shotcrete accelerators before being applied. Fibers are frequently applied in shotcrete to achieve better performance in terms of tensile strength, durability, and antifire function of the lining.
Presupport measures in fractured, yet good rock mass aim to increase standup time. This limits the overbreak, ensuring safe excavation and allowing efficient initial support installation. Spilling (Fig. 4.40) and forepoling are two common presupport measures.
The former is the insertion of tensile strengthening elements in the ground such as dowels. The latter involves the installation of steel sheets or bars. In cases where the rock mass is severely fractured, self-drilling rock reinforcement pipes are used to avoid borehole collapse (U.S. Department of Transportation Federal Highway Administration (FHWA), 2009).
Fig. 3.82 illustrates the excavation process. Rock bolts and prestressed anchors were used as reinforcements; the rock bolts were installed in a pattern of a 2 × 2 m grid over the entire arch and a 3 × 2 m grid on the two sidewalls. Reinforcement in each stage was conducted in the following sequence: (1) install rock bolts, (2) embed prestressed anchors, and (3) place shotcrete. The rock bolt and prestressed anchor were both treated as cable elements with different properties in the numerical model; the shotcrete was modeled as a structural element with a thickness of 0.32 m. A FISH function was written to perform automatic excavation and reinforcement installation.
Table 3.38 lists the input parameters for the structural elements and cable elements in the simulation (Li, 2009). Blasting dramatically affects the rock mass stability. In the simulation, the rock masses within 2 m from the wall of the excavation have been treated as a blast-induced damaged zone with a decreased deformation modulus that is 36% of the original value (Li, 2009).
Table 3.38. Input parameters for the reinforcements (Li, 2009).
In addition to friction at the bolt-rock interface, the bolt also interacts with the rock through mechanical interlock owing to the roughness of the borehole surface. The bolt tube is plastically deformed under the high water pressure during installation to match the rough profile of the borehole surface. A secondary contact stress q2 will be induced at the bolt-rock interface when the bolt tube tends to longitudinally slip along the borehole. Letting i represent the average roughness angle of the borehole surface, the radial contraction of the bolt steel tube, ui, is related to the axial slippage x as follows:
The secondary contact stress is expressed, according to Eq. (4.3.9), as
The magnitude of the secondary contact stress can be demonstrated through the following example. Assume that an inflatable bolt is installed in hard rock. The radius of the borehole is ri = 25 mm, and the roughness angle of the borehole surface is i = 5°. The other relevant parameters are the tube thickness t = 3 mm, and the Young's modulus and the Poisson's ratio of the tube steel are Es = 210 GPa and νs = 0.2, respectively. The secondary contact stress q2 is obtained, according to Eq. (4.3.13), to be 10.5 MPa after 1 mm axial slippage of the bolt. In reality, the secondary contact stress may never reach such a high magnitude before plastic yielding occurs on the tube surface under the asperities of rock. Generally speaking, the secondary contact stress is considerably higher than the primary contact stress in hard rock. In this calculation example, it is assumed that the asperities on the borehole surface are not damaged during bolt slippage. In soft and weak rock, however, the asperities could be easily damaged, and thus, the roughness angle would be significantly degraded after bolt installation. Thus, the secondary contact stress is less important than the primary contact stress in weak rock.
The products under consideration by Huade include FRP rock bolt, vessels, pipelines, duct valves and fittings, cooling tower, spiral, safety helmet, electric fan, cable pipe, railing, grating, crane span, and others. Rock bolt (Fig. 16.18) is widely used in underground mining to provide support to the roof or sides of the cavity. It can be used in any excavation geometry and it is simple and quick to apply, and is relatively inexpensive. The installation can be fully mechanized. The length of the bolts and their spacing can be varied, depending on the reinforcement requirements. However, in aggressive environmental conditions such as in coal mines, steel bolts deteriorate in a matter of days rather than years. FRP rock bolt is particularly suitable in harsh chemical and alkaline environments because of its corrosion resistance. It is durable, lightweight, easy to install, non-conductive, dimension-ally stable under thermal loading, anti-static rating, and can be cut without the danger of sparks. Currently, four production lines are being installed at the Beijing Huade facilities. Furthermore, spiral (Fig. 16.19) is widely used in coal preparation equipment and fine coal washing systems, and is being manufactured using a hand layup method. It is made of FRP composites with a wear-resistant coating.
The directional rock bolts should be designed for tackling loads due to the wheels of the crane on the haunches.
Support must be installed within the stand-up time (Figure 6.1).
While adopting the empirical approaches, it must be ensured that the ratings for the joint sets, joint spacing, rock quality designation (RQD), and so forth are scaled down for the caverns if initial ratings are obtained from the drifts. This is done because a few joint sets and weak intrusions in a drift could be missed. The rock mass quality should be downgraded in the area of a shear zone and a weak zone (see the section Treatment for Tunnels in Chapter 2). A mean value of deformation modulus Em should be substituted for Ed in Eq. (12.1) for estimating the length of wall anchors. Similarly, a mean value of rock mass quality (Qm) and joint roughness number (Jrm) should be used in Eq. (8.9) for assessment of the ultimate support pressure.
Stresses in the shotcrete lining and rock anchors may be reduced significantly by delaying subsequent layers of shotcrete (except initial layers), but no later than the stand-up time. Instrumentation for the measurements of stress and deformation in the roof and the walls of a cavern or in tunnels is a must to ensure a safe support system. Instrumentation would also provide feedback for improvements in the designs of such future projects. Location of instrumentation should be judiciously selected depending upon the weak zones, rock mass quality, and intersection of openings.
Two parallel road tunnels are constructed for six lanes in basalt. The tunnels are D-shaped with diameter (B) of approximately 16 m and with 2 m high side walls with clear spacing of 20 m. The maximum overburden (H) is 165 m. The rock mass parameters are RMR = 73, Q = 10, Ja = 1.0, Jr = 3.0, and Jw = 1.0 (minor seepage from side walls). The construction engineers want a rapid rate of tunneling and life of the support system should be 100 years. The UCS of SFRS is 30 MPa and its flexural strength is 3.7 MPa.
The short-term support pressure in the roof may be assessed by following correlation (Eq. 6.6) for the arch opening, given by Goel and Jethwa (1991):
The ultimate support pressure is read by the chart (Figure 8.2) of Barton et al. (1974) as follows (the dotted line is observed to be more reliable than correlation).
(The rock mass is in non-squeezing ground condition (H < 350 Q1/3) and so f′ = 1.0. The overburden is less than 320 m, so f = 1.0.)
It is proposed to provide the SFRS (and no rock bolts for faster rate of tunneling). The SFRS thickness (tfsc) is given by the following correlation (using Eq. 12.10):
The tensile strength of SFRS is considered to be about one-tenth of its UCS, so its shear strength (qsc) will be approximately 2 × 30/10 = 6.0 MPa, but we will say 5.5 MPa (uniaxial tensile strength is generally less than its flexural strength). Past experience reflects the same information.
The life of SFRS is the same as concrete in a polluted environment of approximately 50 years. Life may be increased to 60 years by providing an extra cover of 5 cm of SFRS. If SFRS is damaged later, the corroded part should be scratched and a new layer of shotcrete should be sprayed that will last for 100 years. For this the recommended thickness of SFRS is tfsc = 13 cm = 21 cm (near portals).
The width of the pillar is more than the sum of the half-widths of adjoining openings in the non-squeezing grounds. The width of the pillar is also more than the total height of the larger of two caverns (18 m); hence the proposed separation of 20 m is safe (Hoek, 2007).
The following precautions need to be taken:
Loose pieces of rocks should be scraped thoroughly before shotcreting for better bonding between the two surfaces.
Unlined drains should be created on both sides of each tunnel to drain out the groundwater and then should be covered by reinforced cement concrete (RCC) slabs for road safety.
Tunnel exits should be decorated with art and arrangements should be made for bright lighting to illuminate the tunnels.
As already noted, the excavated opening is likely to be stabilised by rock bolting with the frequency and length of the rock bolts being determined by the local rock quality. Consideration would have to be given to the choice of materials used in connection with this control method to avoid introducing materials that may be detrimental to repository performance, e.g. organic resins. The spalling of small fragments of rock can be prevented by the use of shotcrete, a surface coating of a specially formulated concrete. Consideration would have to be given to the chemical perturbation that this would cause in the long term, which might necessitate its removal prior to backfilling and sealing the relevant section of the repository. Similar considerations will apply to any use of concrete in structural components of the underground workings, such as floor slabs below waste handling machines, and some programmes are studying the potential development of specially formulated concretes that will minimise the chemical perturbation caused by their use (e.g. Vuorinen et al, 2004).
Groundwater inflows into the excavations can be controlled to some extent by grouting the flowing fractures or more simply by piping the water into a sump if the flow rate does not warrant grouting. The construction of sealing systems to isolate various compartments of the underground system is covered as part of the discussion of engineered barriers in another chapter.
Littlejohn (1993) classified various types of axial failure when using grouted bolts as follows: the bolt, grout, rock, bolt–grout interface, and grout–rock interface. The type of axial failure depended on the properties of individual elements. The shear stress at the bolt–grout interface was greater than at the grout–rock interface because of the smaller effective area. If the grout and rock were of similar strengths, failure could occur at the bolt–grout interface. If the surrounding rock was softer then failure could occur at the grout–rock interface.
Based on pull-out tests of cable bolts in the laboratory and in the field, Hyett et al. (1992) have identified two failure modes in cementitious grouted cable bolt. One mode was radial splitting of the concrete cover surrounding the cable, while the other involved shearing of the cable against the concrete. The former concerns the wedge mechanism but it is rarely observed in the resin grouted bolting system. The shearing mechanism involved crushing of the grouting material ahead of the ribs on the bar, eventually making pull-out along a cylindrical friction surface possible. It should also be noted that as the degree of radial confinement increased, the failure mechanism changed from radial fracturing of the cementitious annulus under low confinement, to shearing of the cement flutes and pull-out along a cylindrical friction surface under high confinement.
Recent research work of failure mode analysis suggests that a cylindrical failure surface around the bolt–resin interface is a predominating failure mode in rockbolting (Cao et al., 2013). It occurs for smooth bars and for very closely spaced rebar bolts (like a screw) along the rib tips of the bar. For rebar bolts, experimental observation suggests that if the embedded length is short and the confining material is stiff, parallel shear failure occurs in laboratory pull-out tests. Fig. 2.27 shows a pull-out test bolt of 75-mm embedded length and confined in 8-mm thick-steel tubes.